A process for the coproduction of zinc and sulphuric acid from zinc sulphide ores

ABSTRACT

A process for the production of zinc and sulphuric acid comprising leaching solid zinc sulphate with an aqueous liquor to form a leach liquor, electrolysing the latter to obtain zinc and a spent electrolyte comprising sulphuric acid, concentrating the sulphuric acid content in the spent electrolyte to precipitate the bulk of the residual zinc sulphate as a solid, separating the latter and so obtaining a sulphuric acid filtrate, and recycling said precipitated zinc sulphate to the electrolysis step in the next cycle by dissolving it in the aqueous leach liquor fed to the electrolysis step. Preferably the starting material (zinc sulphate) is obtained by a controlled oxidation (sulphating) roast of zinc sulphide ore or zinc containing residues obtained in zinc manufacturing processes.

United States Patent [1 1 Bourchier et a1.

[ Aug. 21, 1973 [54] PROCESS FOR THE COPRODUCTION 017 1,443,707 1/ 1923 Gepp 75/ 120 X ZINC AND SULPHURIC ACID FROM ZINC $=PP g:

, oung SULPHIDE ORES 2,435,340 2/1948 Christensen 23/125 X [75] Inventors: Kenneth Arthur Bourchier, 2,754,174 7/1956 Roberts 204/119 X Bundoora, Victoria; Donald 3,434,947 3/1969 Steintveit 75/120 X Fergumn Stewart, Doncaster, 3,493,365 2/1970 Pickering et a1 75/120 Victoria, both of Australia Primary Examiner-0scar R. Vertiz [73] Asslgnee: Imperial Chemical Industries of Assistant Examiner chafles Rodman f and zealand Limited Attorney-Cushman, Darby and Cushman Victoria, Australia [22] Filed: Aug. 4, 1970 [57] ABSTRACT [2]] App! NM 60 908 A process for the production of zinc and sulphuric acid comprising leaching solid zinc sulphate with an aqueous liquor to form a leach liquor, electrolysing the latpp Priority Data ter to obtain zinc and a spent electrolyte comprising Aug. 20, 1969 Australian" 59791/69 sulphuric acid, concentrating the sulphuric acid con- May 21, 1970 Australia 1283/70 tent in the spent electrolyte to precipitate the bulk of the residual zinc sulphate as a solid, separating the lat- [52] 0.8. CI 75/115, 75/120, 423/95, ter and so btaining a sulphur a id fil rat a r y- 423/106, 423/530, 204/119 cling said precipitated zinc sulphate to the electrolysis [51] int. Cl C0lg 9/00, COlb 17/72 step in the next cycle by dissolving it in the aqueous [58] Field of Search 23/125, 167; 75/120; leach liquor fed to the electrolysis step. Preferably the 204/119; 423/53, 106 starting material (zinc sulphate) is obtained by a controlled oxidation (sulphating) roast of zinc sulphide ore [56] References Cited or zinc containing residues obtained in zinc manufac- UNITED STATES PATENTS luring Processes- 1,241,966 10/1917 Hanley 23/125 x 4 Claims, 1 Drawing Figure AIR ORE so H suwnumc H2 504 ROAST CONCEQNI3- EPRODUCTD SPENT ELECTROLYTE CALCINE LIQUOR ZINC SULPHATE CRYSTALS 2 4 LEACH ELECTROLYSIS figl stats EZZ'SE" FILTRATE SEPARATION LEACH PURIFICATION SECONDARY SOLUTION leagtfiosucr) RESIDUE EAD SULPHATE ENRICHMENT CONCENTRATE OF E LEAD SULPHATE PRODUCT PATENTEB M1921 I875 sivsalssz AIR one so; H 0

SULPHURIC 5 H so ROAST A 4 CONCENTRAT- PRODUCT) SPENT ELECTROLYTE CALCINE L'QUOR zmc SULPHATE CRYSTALS zmc c LEA H WELECTROLYSIS fi o PURIFIED LEACH LEACH SLURRY FILTRATE LEACH SECONDARY SEPARATION SOLUTION PURIFICATION 'aa paoouc'r) METALS RESIDUE LEAD SULPHATE ENRICHMENT A PROCESS FOR THE COPRODUCTION OF ZINC AND SULPHURIC ACID FROM ZINC SULPIIIDE ORES Thus invention relates to the recovery of zinc values from zinc sulphate solutions; in particular it relates to the recovery of zine values from zinc sulphate solutions obtained from zinc bearing materials, for example, from sulphide ores such as mixed zinc sulphide, lead sulphide, iron sulphide and silver sulphide ores or from zinc containing residues obtained in zinc manufacturing processes such as an electrolytic manufacturing process.

The conventional process of winning zinc from zinc sulphide ores is exhaustive oxidation; the sulphide is roasted to break the zinc-sulphur bond, zinc is converted to zinc oxide, sulphur is oxidised to sulphur dioxide which, in turn, is catalytically oxidised further to sulphur trioxide and converted to sulphuric acid; the zinc oxide is dissolved in spent recycle electrolyte liquor (electrolysis effluent) which contains sufficient weak sulphuric acid for this purpose; then, after removal of by-product metals and impurities, this solution is electrolysed to produce the desired metallic zinc and said spend recycle electrolyte liquor (electrolysis effluent).

In this process complex, expensive and vast plant capacity is required for removing sulphur from zinc sulphide as sulphur dioxide, for purifying the sulphur dioxide, for oxidising this sulphur dioxide to sulphur trioxide, and for combining the latter with water to form sulphuric acid. In addition it is necessary that this plant capacity be located at the site at which the ore is roasted.

We have now found a process, by which the energy expended in the conventional electrolysis process to form zinc is utilised to produce simultaneously sulphuric acid as a product available for sale and the need to generate an equivalent amount of sulphur dioxide, purify it, oxidise and convert it to sulphuric acid and the cost of the associated plant can be avoided. In addition we have evolved a workup process in which the manufacture of sulphuric acid from zinc sulphide is not physically linked to the roasting process and which, therefore, permits greater flexibility in plant layout and geographic location of the several process steps.

Accordingly we provide a process of recovering zinc values from a zinc sulphate solution, which comprises:

A. electrolysing said solution to yield metallic zinc as a product and spent electrolyte liquor as an intermediate, which is an aqueous solution of sulphuric acid and zinc sulphate;

B. concentrating the sulphuric acid in said spent electrolyte liquor so that its liquid phase contains at least 600 g/l preferably between 900 and 1,150 g/l of sulphuric acid and the bulk of the zinc sulphate is precipitated, separting this precipitate from the liquor and so obtaining sulphuric acid as a product; and

C. redissolving said precipitate of zinc sulphate in the feed liquor to the electrolysis step A and thereby fortifying it so as to obtain from it additional zinc as the product.

This process has several important advantages over the conventional process.

In conventional electrolysis processes, the spent electrolyte liquor (electrolysis effluent) is recycled for two reasons:

first: to provide an acid to dissolve the oxide calcine and second: to return to the system the large proportion of the zinc sulphate which it still contains and which would otherwise be lost in the liquor.

In our process, neither of these reasons pertains. Firstly, as already stated, sulphuric acid is not required to dissolve the calcine obtained from the roasting process; instead virtually all of the sulphuric acid formed in the electrolysis step is available for conversion into a saleable product. This represents superior utilisation of electric energy, since no additional power is required to obtain two (Zn and sulphuric acid) instead of one (Zn) electrolysis product. Consequently sulphuric acid plant capacity associated with the formation, the oxidation andthe absorption of the bulk of the sulphur dioxide derived from zinc sulphide is not required. The second reason for recycling spent electrolyte liquor the need to recover zinc pertains in the conventional process even if the calcining process is modified to yield zinc sulphate which can be dissolved without the use of acid.

In our process this second reason, too, has been obviated by a solids-recovery step.

The virtual absence of recycles of liquors in our process has further advantages; the conventional recycle liquors contain soluble impurities: which, on repeated recycling, build up to undesirable levels; consequently additional purification steps have to be introduced. In the process of the present invention the recycled zinc sulphate is crystallised from sulphuric acid as a solid;

this solid is relatively pure and, but for the small amount of sulphuric acid which is bound to be associated with it, carries with it very small amounts of impurities only. Consequently less purification is required. The purification step is less complex, too, because the leach liquor is less acidic, hence less neutralising is necessary before purification.

Yet another advantage of our process is its flexibility; since the sulphuric acid is derived from the electrolyte and not the roasting step, its manufacture is not of necessity tied to the siteof roasting; electrolysis, our zinc recovery and our sulphuric acid manufacture may be carried out in geographic isolation from the site of roasting; this alleviates the problems associated with environmental pollution by sulphur dioxide production and with transport of sulphuric acid. Greater flexibility results also in the electrolysis step; in the prior art processes the upper limit of the zinc sulphate concentration in the cell feed liquor (electrolyte) was given by the amount of weak sulphuric acid which can economically be generated in the spent electrolyte; using the recycle of solid zinc sulphate according to the present process and leaching of solid zinc sulphate to make up the fresh electrolyte (cell feed), higher concentrations of zinc sulphate and operation over a greater range of concentrations can be attained. This permits flexibility, a higher degree of decomposition per pass through the electrolytic cell, better energy efficiencies (lower operating voltage at higher electrolyte concentration), higher capacity per floor area and greater current density.

The conventional electrolysis of zinc sulphate uses a zinc oxide calcine as its raw material, that is electrowinning of zinc is combined with a conventional oxidising roast producing zinc oxide and sulphur dioxide from zinc sulphide. This combinedprocess has several disadvantages; firstly very large amounts of sulphur dioxide are formed which not only require large plant for conversion to sulphuric acid, as already discussed, but which, in such quantities, present problems of environmental control air pollution with damage to plants and habitation and which consequently restrict the choice of geographic location of the roasting plant. In addition the crude zinc calcine, mainly zinc oxide mixed with gangue, is far less readily dissolved in weak sulphuric acid (spend recycle electrolyte liquor) than would be expected from the behaviour of pure zinc oxide. Considerable excess acidity must be provided to achieve practical rates and efficiencies of leaching; this acidity must either subsequently be neutralised with zinc oxide or zinc, or the leaching must be conducted in two or more stages with decreasing acidity; the former reduces the net plant output, the latter requires additional plant; both are costly.

We have now recognised that the ability of our process to utilise zinc sulphate as its starting material enables us to devise a new combination process.

Accordingly we provide a process of recovering zinc from an ore comprising zinc sulphide which process comprises, in combination, converting the zinc sulphide into an aqueous zinc sulphate solution, and then electrolysing said solution to obtain zinc and spent electrolyte liquor, concentrating the sulphuric acid in the spent electrolyte liquor to obtain sulphuric acid and a precipitate of solid zinc sulphate and redissolving said zinc sulphate in the feed liquor to the electrolysis step so as to obtain additional zinc as hereinbefore described.

The zinc sulphate may be obtained by several methods. It may be prepared by dissolving zinc oxide (e.g. the zinc oxide contained in conventional calcine) in the sulphuric acid concentrate obtained in our concentration step B from spent electrolyte; this departure from the conventional process of dissolving calcine in weak spent electrolyte can, at times, be useful for dissolving particularly intractable calcines, or parts thereof.

Alternatively, zinc sulphate could be obtained by a pressure leaching process such as described by V.N. Mackiw and H. Veltman in the Canadian Mining and Metallurgical Bulletin for January, 1967, pp. 80 85.

Finally, and this is a preferred embodiment of our invention, zinc sulphate is obtained by applying a process which may be called sulphating roasting to zinc sulphide ores. By sulphating roasting we mean a process of roasting zinc (or lead) sulphide ore under conditions of controlled oxidation such that, in the overall effect, the sulphidic sulphur remains bound to zinc (or lead), the bulk of the zinc (or lead) sulphide is oxidised to sulphate and the formation of sulphur dioxide from zinc or lead sulphide is minimised or substantially eliminated. Sulphating roasting, as far as we are aware, has not hitherto been used technically in the winning of zinc from zinc sulphide; the application of a sulphating roasting process of zinc sulphide ore and the combination of this process with our above described sequence of leaching, electrolysing, concentrating and precipitation provides an entirely new process sequence of winning zinc from zinc sulphide ores which permits the selective recovery of zinc from crude ore, produces sulphuric acid without the generation of sulphur dioxide from the bulk of the zinc sulphide, obviates the formation of zinc oxide and overcomes the difficulties associated with dissolving it.

Accordingly we provide a process which comprises, in combination:

l. roasting an ore comprising zinc sulphide at a temperature in the range of 500 to'800C in an atmosphere which contains oxygen and sulphur dioxide to convert the ore to a calcine, comprising the bulk of the zinc originally present in the ore in the form of zinc sulphate;

2. leaching said calcine with water or an aqueous solution at a temperature in the range 10 to 100C, using a solidszliquid ratio up to 2:1 to yield a leach slurry;

3. subjecting said leach slurry to a solids/liquid separation process to yield a leach solution which comprises as dissolved zinc sulphate at least percent of the zinc contained in said calcine, and a residue, which contains silicaceous gangue, insoluble iron and lead compounds;

4. electrolysing said leach solution, optionally and preferably fortified by dissolving in it solid recycle of zinc sulphate from step 5 to yield, as a product, metallic zinc and, as an intermediate, spent electrolyte liquor, which is an aqueous solution of sulphuric acid and zinc sulphate;

5. concentrating sulphuric acid in said spent electrolyte liquor so that its liquid phase contains at least 600 g/l and preferably between 900 and l,150 g/l of M and the bulk of the zinc sulphate is precipitated, separating this precipitate from the liquor and so obtaining sulphuric acid as a product. As described above, preferably the precipitated zinc sulphate is redissolved in the feed liquor to the electrolysis step and recycled in subsequent operations.

This combined process is particularly useful for certain intractable Zn/Pb ores which are not readily amenable to the conventional processes of oxidative roasting. Ores of this type are characterised by an extremely tine-grained sulphide mineralisation in a siliceousdolomite matrix, typically represented by the ore body known as the McArthur River deposit of the Northern Territory of Australia. This ore is the subject of a petrological report A Mineralogical Examination of the McArthur Lead-Zinc-Silver Deposit by N.J.W. Croxford, presented at the Australasian Institute of Mining and Metallurgy Conference, Mount Isa, 1967. Ores of this type, as well as posing difficulties in primary metallurgical processing, present problems in the subsequent processes of extraction of their useful metal values. They can usually be concentrated by flotation, although with difficulty, but no economical separation of the two major metal values, lead and zinc, has hitherto been achieved. If pyrometallurgical extraction processes are applied to the ores or their concentrates to extract the lead and zinc values by smelting, a slag is formed of undesirable viscosity, the separation efficiencies are low and large amounts of sulphur dioxide are formed. Alternatively, hydrometallurgical processes may be applied to the ore or concentrate, or to the calcined products derived from them, to product solutions of zinc salts suitable for electrolysis. The lead values so obtained are generally insoluble, hence a separation of lead and zinc can be achieved. However, on leaching with spent electrolyte, the iron present in the ores and concentrates is solubilised along with the zinc and presents a major problem in the subsequent electrolysis step which is used to obtain metallic zinc. As a consequence the McArthur River deposits, although a major Australian mineral find and subject of much investigation over the years, to-date have not yet been exploited. These difficulties can be overcome by the present process', we have shown that the ore can be resolved with high yields, that by-product metals such as lead, silver, cadmium and copper are readily recovered, and that iron, the separation of which is particularly troublesome, and other impurities are readily removed. The removal of iron is particularly effective when the controlled oxidation technique is used, by which iron sulphide is oxidised preferentially to its oxides, while zinc and lead sulphide are oxidised selectively to the respective sulphates.

Accordingly a particularly preferred process of our invention is characterised in that zinc sulphide ore or ore concentrate is roasted in two steps, wherein the first step comprises roasting said ore in the presence of oxygen, preferably air, and sulphur dioxide at a temperature between 50 and 800C, preferably 650 and 750C, to convert the bulk of the zinc sulphide into zinc sulphate and the second step comprises maintaining the roast at a temperature between 400 and 700C, preferably between 450 and 600C in an atmosphere from which sulphur dioxide has been substantially displaced by air or oxygen or purged to decompose water-soluble iron compounds.

The subsequent extraction, purification, electrolysis, precipitation and concentration steps are then carried out as above described.

The individual steps of the present process are described below;

I. Roasting The invention is applicable to ores containing zinc sulphide which may or may not be partially oxidised. it is preferred to apply the process of the invention to the concentrate of such ores as are obtained by conventional concentration processes.

If an ore is used without concentration, the preferred particle size range is -60 to +300 B.S.S. mesh. Concentrate is suitable in the particle size range obtained from the concentrators.

As already stated, the roasting process is preferably carried out that is in two steps, the first involving the formation of sulphates and the second involving the decomposition of iron sulphate. Alternatively the roasting may be carried out in one step at a temperature between 600 and 750C.

The suitable apparent ratio of S0 0 present during the initial roasting stage is between and 1:2 and the preferred ratio is between 5:1 and 5:6. It is referred to as apparent ratio" because it is known that some of the sulphur dioxide present is oxidised to sulphur trioxide; these ratios are therefore defined as the amount of sulphur dioxide plus sulphur trioxide (expressed as sulphur dioxide) present in relation to the amount of oxy ing solution is in the range 5.0 to 5.5 to lower the amount of iron dissolved. The preferred base is ammonia and the preferred temperature range is 60 to C. The preferred solid/liquid ratio is in the range from 1:4 to 1:1. By solid/liquid ratio we mean the weight ratio of total, dissolved and undissolved, solids to water.

3. Phase Separation and Purification The leach slurry, essentially, consists of a water phase comprising in solution, zinc sulphate, some iron sulphate, and soluble salts of other trace metals associated with the ore, and a solid phase comprising lead sulphate, iron oxides, ferrities, silicaceous gangue and insoluble compounds of other trace metals associated with the ore.

These phases are separated by conventional pr0 cesses involving the use of hydrocyclones, and/or thickeners and/or filters.

The liquid pahse, hereinafter referred to as leach solution, may and usually does contain impurities, such as iron, arsenic, cadmium, antimony and copper, which, if present in sufficient quantities, are detrimental to the subsequent electrolysis step, and must therefore be removed before electrolysis. The methods of keeping these impurities below the levels acceptable for electrolysis are known from the prior art. One widely practised method is to add zinc dust to the leach solution; a zinc dust precipitate forms, is separated from the solution and the trace metal values contained in it may be reqovered, if desired. Further purification steps may be necessary depending on the character of the ore bein treated. 1

As stated, it is preferred to fortify the leach solution after purification with the solid ziinc sulphate, which is obtained in the concentration step from the concentrated spent electrolyte liquor. The foritifed leach solution is then subjected to electrolysis. The solid residue from the leaching step comprising gangue and usually lead is worked up as described in step 5 4. Electrolysis The electrolysis step is carried out in a manner known e.g. from Mantel], Industrial Electrochemistry, McGraw Hill, 1950 Edition, pp. 362 to 364, with the following important differences:

a. the feed electrolyte, preferably, is fortified as described, with solid zinc sulphate, which may constitute up to one third of the total zinc content in the feed electrolyte; consequently the degree of electrolytic decomposition per pass may be higher than in the prior art;

b. the spent electrolyte is not recycled to the extraction step but worked up as described below. From the electrolysis metallic zinc is then obtained in the conventional manner.

5. Concentration The limit to which electrolysis can be driven depends on the amount of sulphuric acid formed and the residual amount of zinc dissolved in the spent electrolyte. Complete electrolysis is not possible; consequently some zinc will remain in solution and must be recovered. To this end the sulphuric acid in the spent electrolyte liquor is concentrated until the concentration of free sulphuric acid is greater than 600 g/l, preferably between 900 g/l and 1,150 g/l. At a concentration of 900 g/l substantially all of the zinc will have crystallised out as zinc sulphate, which is separated from the sulphuric acid, e.g. by decantation, filtration or preferably by centrifuging.

By concentrating the sulphuric acid we mean increasing the sulphuric acid content of the spent electrolyte liquor to the desired level by one of several methods; these are:

i. concentrating the spent electrolyte liquor by evaporating water; or,

ii. adding more concentrated sulphuric acid (e.g.

98% H 80 to the spent electrolyte liquor; or,

iii. absorbing sulphur trioxide in the spent electrolyte liquor; or,

iv. a combination of steps i and ii; or,

v. a combination of steps i and iii; or,

vi. a combination of steps ii and iii; or,

vii. a combination of steps i, ii and iii. The combinations of steps iv and vii inclusive, particularly the use of more concentrated sulphuric acid or sulphur trioxide after the evaporation, have the advantage that less heat is required than when evaporation alone is used. The more concentrated sulphuric acid may be obtained from the sulphuric acid produced in the system of absorption of sulphur trioxide or, of course, from an outside source, e.g. purchase.

The mother liquor from the concentration/filtration squence is sulphuric acid, as useful byproduct. If it is required in a more concentrated form, then the abovementioned concentration step may be extended by addition of sulphur trioxide to the concentrated electrolyte liquor in an absorption step such as is normally used in sulphuric acid manufacture; alternatively, it may be concentrated further by evaporation.

The zinc sulphate, which may contain some entrained sulphuric acid, is preferably recycled to the leach solution, most preferably between the purification and electrolysis steps, as described above and as indicated in the diagram of the FIGURE.

6. Work-up of Solids after Leaching The solid phase or residue from the leaching step,

commonly contains lead in the form of sulphate in quantities worthy of recovery. The lead values may be recovered from the leach residue by conventional pro cesses, e.g. smelting. Preferably, however, the lead sulphate from the residue is first concentrated by flotation; as a concentrate it is suitable for further processing to recover the lead values. This concentration process may be enhanced by treating the residue with sulphuric acid, preferably from the processing of the spent electrolyte liquor, to remove most of the iron content leaving a residue of lead sulphate and silicaceous gangue.

Alternatively the lead may be concentrated by extraction with sodium hydroxide or by other methods. 7. Combined Process The combined process of this invention is illustrated in the diagram of the FIGURE of this specification in which each numbered box represents a step as described. The raw materials are the ore or ore concentrate, air, water and sulphur dioxide and the effluents and products areas labelled. The sulphur dioxide used in the process may be generated within the process from the sulphur associated with iron and possibly with the lead. The primary products are zinc metal, lead sulphate concentrate and sulphuric acid. Other valuable secondary products such as copper and silver may be obtained.

Our invention may also be used to extract the zinc values from normally intractable residues obtained in zinc manufacturing processes such as an electrolytic manufacturing process.

Accordingly we provide a process which comprises, in combination:

1. roasting a zinc residue at a temperature in the range of 500 to 800C in an atmosphere which contains oxygen and sulphur dioxide to convert the residue to a calcine, comprising the bulk of the zinc originally present in the residue in the form of zinc sulphate;

2. leaching said calcine with water or an aqueous solution at a temperature in the range 10 to 100C, using a solidszliquid ratio up to 2:1 to yield a leach slurry;

3. subjecting said leach slurry to a solids/liquid separation process to yield a leach solution which comprises as dissolved zinc sulphate at least percent of the zinc contained in said calcine, and a residue, which contains silicaceous gangue, insoluble iron and lead compounds;

4. electrolysing said leach solution, optionally and preferably fortified by dissolving in it solid recycle of zinc sulphate from step 5 to yield, as a product, metallic zinc and, as an intermediate, spent electrolyte liquor, which is an aqueous solution of sulphuric acid and zinc sulphate;

5. concentrating sulphuric acid in said spent electrolyte liquor so that its liquid phase contains at least 600 g/l and preferably between 900 and 1,150 g/l of 1-1 and the bulk of the zinc sulphate is precipitated, separating this precipitate from the liquor and so obtaining sulphuric acid as a product. As described above, preferably the precipitated zinc sulphate is redissolved in the feed liquid to the electrolysis step and recycled in subsequent operations.

The leach solution may optionally be treated according to processes known in the art to remove dissolved iron. For example, air may be bubbled through the solution to precipitate the iron as iron hydroxide which may subsequently be removed. Another suitable method is to treat the solution with a mixture of ammonium or sodium salt and manganese dioxide and to remove the precipitated iron compounds.

ln our invention any suitable method of separating the bulk of the zinc sulphate from the liquor may be employed. Suitable methods known in the art are, for example, salting or freezing out the zinc sulphate or the use of various known dialysis procedures using semipermeable membranes. Preferably we use the concentration procedure described hereinabove.

Accordingly we provide a process of recovering zinc values from a zinc sulphate solution which comprises, in combination:

a. electrolysing said solution to yield metallic zinc as a product and spent electrolyte liquor as an intermediate which is an aqueous solution of sulphuric acid and zinc sulphate;

b. separting the zinc sulphate from the said spent electrolyte liquor; and

c. redissolving said zinc sulphate in the feed liquor to the electrolysis step aand thereby fortifying it so as to obtain from it additional zinc as the product.

The zinc sulphate solution may be derived from any zinc bearing material. The invention is especially useful when conventional methods of winning zinc from the zinc bearing material have hitherto been considered to be difficult or uneconomic.

The zinc sulphate solutions, may for example, be derived from a zinc bearing material such as an ore or a Example 23 An aqueous solution of sulphuric acid and zinc sulphate of the following composition:

zinc residue obtained in a zinc manufacturing process. il Znso4 50 g Sulphunc acld H,SO. 98 g /1 Preferably the ore comprtses zinc sulphtde. In th1s specifi ation th t i c e 'T ore 8 used to mean me come wh1ch 1s typlcal of spent electrolyte hquor, was concentrate or pellettzed ore concentrate.

The SUI hatin mast hereinbefore described ma be trated by evaporatton. The amounts of zmc dissolved ca rie 3; using e e a fixed bed or n d and the amount of zinc sulphate precipitated at various o ore pa c e g a 8 stages of concentration were determined. The results are iven in Table 3. The chotce of the type of bed 1s related to the stze of g the operation. Thus on a small scale it may be economi- TABLE 3 cally desirable to use a fixed bed, whereas fluidized beds may become more convenient as the scale of op- .15 Zinc Sulphate eration is increased. precipimed The invention is now illustrated by, but not limited w fj flififi gif to, the following examples. H so I of in mining Example 1 to 14 inclusive fi if f Zn 2 501mm Sampels of a flotation concentrate of McArthur River ore of average analysis of 27.0% Zn, 12.8% Fe, 28 A88 50 50 0 11.0% Pb and 29.1% s were roasted under the condi- 3 245 B 3 5'? tions outlined in Table l. Aftercompletion of the roast- 600 163 47.4 7.7 8415 ing step the calcines were extracted in boiling water (5 588 {$3 5 6-5 32- g calcine in 50 ml water). The amounts of zinc and iron 900 6 :2; extracted thereby were determined by X-ray fluores- :988 33.1 1.; 0.147 99.71 t. 0.134 99.73 cence and atomic absorpt1on techniques. The results 1200 816 L2 0098 99.80 are given 1n Table 1.

TABLE 1 Roasting 1st stage Roasting 2nd stage Percent of P 1 Temp., AirISOz Time, Temp., Time, Zn in ore fi iri o e 0. ratio hrs. C. AirzSOrratio hrs. extracted extracted 50o 3.33:1 4 500 1 29.5 as 050 3. 33:1 4 600 1 s5. 3 4. 5 700 3. 33:1 3 600 1 s0. 7 3. 25 700 0.75:1 2 600 1 46.2 0.0 700 1. 08:1 2 600 1 91. 5 11. 4 700 20:1 2 600 1 87.8 3.4 700 25:1 2 600 1 91.8 6.9 700 3. 33:1 3 600 1 86.2 2. 9 700 :1 2 600 1 37.4 2.7 700 3.0:1 2 600 1 88.0 2.3 700 7.5:1 2 600 1 75.0 0.0 700 2. 5:1 2 600 1 91. 2 10. 0 700 3.33:1 3 600 3.33:1 1 90.5 8.1 700 2. 5:1 2 600 Flow of air and SO; discontinued. 1 92. 7 9. 7

Example 15 to 22 inclusive A calcinefrom a sample of McArthur River ore concentrate was prepared by three hours roasting at 700C under an airzSO atmosphere of 3.33:1, followed by treatment at 600C for 1 hour with air only input. The calcine was leached with the solution shown in Table 2 at a solids/liquid ratio of 1:10 and at a temperature of 80C.

The proportions of zinc and iron originally present in the calcine and leached out by the treatment are reported in Table 2. It is seen from this that if the pH of the leach liquor is less than 1.0, inacceptably large amounts of iron are dissolved.

.-'Ih0 nirzSOn ratio defines the atmosphere in the furnace under equilibrium conditions.

Example 24 All analysis figures relevant to this example are shown in Table 4. 400.0 g of calcine (1) from a sulphating roast carried out as described in Example 13 was leached with 800 ml of distilled water at C for approximately one-half hour with continuous stirring. The residue (11) was filtered from the slurry, washed with water, and dried. This material was analysed by x-ray fluroescence. The filtrate or leach solution (111) was diluted to one litre with distilled water and 50 ml. of this solution was retained for analysis by atomic absorption spectroscopy.

950 ml. of leach solution was treated with 20 g. of coarse zinc dust while stirring and heating the solution. The black solid residue from this treatment was filtered off, ignited and weighed (22 g.). This material was dissolved in nitric acid and, on analysis by atomic absorption spectroscopy, was found to contain 0.73 percent copper and 0.07 percent silver and 84 percent Zn. The leach solution was further treated by bubbling air through the solution and heating. The leach solution was filtered to remove the ferric hydroxide precipitated during this procedure and 50 ml of the thus purified filtered solution (IV) taken for analysis by atomic absorp- 1 l 12 tion spectroscopy. Zinc sulphate (116 g Zn SO H O) Example 26 (V), the expected yield from the above described con- 1 litre of electrolyte liquor at a concentration of 40.6 centration step (Example 23) applied to the spent elecg/l zinc as zinc sulphate and 93 g/l with respect to sultrolyte liquor was dissolved in the remaining 850 ml. of phuric acid was prepared as described in Example 25. each solution and the resulting solution (Vi) electro- 5 This solution was concentrated in a stepwise manner,

lysed in a cell with a lead anode and an aluminum cathzinc sulphate monohydrate being filtered off at various ode. Electrolysis was carried out for M5 hours at 4, sulphuric acid concentrations, washed with water and volts and 3.8 amps in a cell with a cathode area of 25 d i d, and the washings returned to the liquor being il'lS. At the end Of thiS period the zinc (VII) deposited concent ated In'this manner 102 g of zinc ulphate was pp from the cathode Washed, dried and monohydrate was precipitated and separated from the Weighed- The ac d Solution remaining in the electrolysis solution. 49. ml of sulphuric acid at a concentration of ell plus w g p electrolyte liquor had 1,799 g/l was added to the 100 ml of the filtrate having 8 Volume of 1.200 ml 50 ml of this solution was take a concentration of 795 g/l of sulphuric acid to produce for analysis. 147.2 ml of final acid at a concentration of 1,114 g/] The remaining 1,150 ml of the spent electrolyte with respect to sulphuric acid and 1.16 g/l with respect q was concentrated until it contained 1,030 g to dissolved zinc. This addition of acid caused the prez J During this concentration Zihc sulphate cipitation of a further 4.4 g of zinc sulphate monohymonohydrate (IX) (identified by X-ray diffraction) was dram precipitated. The electrolysis step had removed more Example 27 zinc than was expected and hence after cooling to C 20 All l i figures relevant to hi example are a total of only 105 g of ZnSO -H O precipitated and shown in Table was separated from the sulphuric field Product by This example shows the application of the process to filtration. Zinc balances across the various steps are rethe normally intractable residue obtained i a convenported in'Table 4.

TABLE 4.ANALYTICAL DATA AND ZINC BALANCE Solids on w./w. basis,

liquids w./v. Zinc balance as percent 0! Zn in- Eleetrol- Spent Leach ysis electro- Item 0! Ex. 24 Sample Percent Zn Percent Pb Percent Fe Calcine solution solution lyte I Sulphated calcine II Leach residue III Leach solution VIII S ent electrolyte liquor IX Z nc sulphate preci itate X Sulphuric acid pro uct Example 25 tional electrolytic zinc plant. The residue used con- All analysis figures and zinc balances relevant to this Sisted i l f i ferrit d contained 232% Example are shown in Table 5. zinc, 31.8% w/w iron and 0.24% w/w Cu.

Cond tions for leach ng and solution p rific n The zinc plant residue was subjected to a sulphating were as in Example 24 with the exception that the leach roast ith a fi t ta f 4 h urs at a t rat r f solution used contained 0.48 g of ammo a a d further 700C and an air to sulphur dioxide ratio in the roaster ammonia was added during iron removal in order to gases of 8:1 and a second stage roast at 600C for onebring the solution pH up to 5.5. Zinc sulpha m n yhalf hour with air only in the roaster gases. 600 g of the drate prepared in Example 24 was added to h leach calcine (1 of Table 6) from this roast was leached with solution prior to electrolysis. Electrolysis was carried 800 l of distilled water at 80C for approximately out at 4 volts and 3.85 amp for a period Of 16.7 h u one-half hour with continuous stirring. The residue (2) in a cell with a cathode area of 12 ins. Concentration was filt d fr m th sl rr wa h d with wat r d of the acid solution after electrolysis was continued undried. This material was analysed by X-ray fluorestilit contained 956 g H,SO /litre. Otherwise procedures rr spectroscopy, Th filtrat r l a h s l ti n (3) were as described in Example 24. was diluted to 1 litre with distilled water and 50 ml of TABLE 5.-ANALYTIC'AL DATA AND ZINC BALANCE Solids on w./w. basis,

liquids w./v. Zinc balance as percent 01 Zn in Electrol- Spent Leach ysis olectro Item 0! Ex. 26 Sample Percent Zn Percent Pb Percent Fe Calclnc solution solution iyto i Sulphated caiclne 21. 6 8.4 Leach residue 3.6 20.6 Leach solution 8.0 0.004 V- Purified leach solution 7. 6 ND V- Zinc sulphate recycle precipitate ND ND VI- Electrolysis liquor 7. 94 ND V Zinc metal 99 ND VIII Spent electrolyte liquor 2. 38 ND IX Zinc sulphate preci itate 16.1 ND X Sulphuric acid pro uct 0.61

In this example zinc compounds were co-precipitated with the iron oxides during the purification stage.

13 14 this solution was retained for analysis by atomic ab- We claim ri spectroscopy. 1. A process which comprises in combination:

in order to reduce the iron content ofthe leach soluaroasting a Zinc bearing material selected from th tion, 950 ml of leach solution was treated with 7 g of group consisting of iron-containin zinc sulphide ammonium sulphate and l g of manganese dioxide. The ore and iron-containing residues from a zinc manusolution was boiled for 4 hours and allowed to stand facturing process at a temperature in the range overnight before filtering. The precipitate so produced 500C to 800C in an atmosphere which contains weighed 75 g and on analysis by X-ray fluorescence oxygen and sulphur dioxide to convert the zinc spectroscopy was found to contain 49.0% w/w iron and bearing material to a calcine comprising the bulk 5.7% w/w zinc. The filtrate was made up to 1 litre and of the zinc originally present in the zinc bearing was further purified by addition of 20 g of coarse zinc material in the form of zinc sulphate;

dust while heating and stirring the solution. The residue 1 l hin s id l i ith wagerm an aqueous from this treatment was filtered washed, dried and i at a [empgfature i h range f 10 to weighed (31.4 g). This material was dissolved in nitric 100 using a i liquid ration up to 2; to yield acid and, on analysis by atomic absorption spectrosa m slurry;

PY, was found to contain 23% pp and 61% c. subjecting said leach to a solids/liquid separation w/w zinc. The leach solution was made up to 1 litre and process to yield a [each solution which comprises 50 ml of this S taken for analysis y atomic as dissolved zinc sulphate at least 70 percent of the abSQTPhOh p py- To the remaining 950 ml of zinc contained in said calcine, and a residue, which Solution was added 137 8 of Zinc sulphate y' contains silicaceous gangue, and any insoluble iron drate (5), the solution was made up to l litre an 50 ml and had compounds;

f this Solution taken l/ The remaining d. electrolysing said leach solution to yield, as a prod- 950 ml was electrolysed for 17 hours at 4 volts and 5 ct. mcmmc zinc and, as an intermediate spcm amps in a cell with a cathode area of ins. At the end d w- 1 li which i an aqueous solution f of this time the zinc deposited (7) was stripped from 25 l h ri a id and zinc sulphate;

the cathode, Washfld dried and welghed- The acid Sollle. concentrating sulphuric acid in said spent electrotion remaining in the electrolysis cell plus washings l t li r so th t it li id Phase tai t least p electrolyte liquor had a ume o 1.2 600 g/l of mso and the bulk of the zinc sulphate ml, 50 ml of this solution was taken for analysis. i r i itated; d

The remaining 1,150 mi of the spent electrolyte lif. separating this precipitate from the liquid and so quor was concentrated until it contained 1,127 g H,SO, obtaining sulphuric acid as a product, the precipi- /litre. During this concentration zinc sulphate monohytated zinc sulphate being redissolved in the feed lidrate (9) was precipitated and separated from the sulquor to the electrolysis step and recycled in subsephuric acid product (10) by filtration. quent operations,

The numerals l to 10 inclusive refer to Table 6. said zinc bearing material being roasted in step (a) in TABLE 6.-.ANALYTIOAL DATA AND ZINC BALANCE Zinc balance as percent of Zn-in- Solids on w./w. basis,

liquids w./v. Electrol- Spent Item of Leach I ysis electro- Ex. 27 Sample Percent Zn Percent Fe Calcine solution solution lyto 1 Sulphated calclne Leach residue....

. Leach solution.

. Purified leach solution.

Zinc sulphate 0 Elsetrolyslsllqunr Z. Zlncmotal s I. Hpcnl. olnctrnlytu llquor Z nc sulphutcdprocl aim... to Sulphur cool pr 0 not Example28 two steps, wherein a first step comprises roasting said 100 ml of electrolyte liquor at a concentration of 30 matel'ial in an atmosphere comprising oxygen and sul g/l zinc as zinc sulphate and 100 g/] with respect to sul- P dioxide at a temperature in the ge f om 500 h -i i was prepared as described in Example 25 to 800C to convert the bulk of the zinc into zinc sul- This solution was treated with 365 ml of 25% w/w Pilate and the Second p comprises maintaining the oleum. The heat of mixing caused some of the liquor to mast at a temperature in the range from to 700C evaporate and the final solution contained 915 g/l of in 'h f from which Sulphur dioxide has been sulphuric acid. The precipitated zinc sulphate monohyf f displaced by a member of the group drate was filt d off d ft drying weighed 55 sisting of air an oxygen to decompose water-soluble Example 29 "On compounds.

100 ml of electrolyte liquor at a concentration of 30 T. mull-ding to claim 1 wherein he W i g/l zinc as zinc sulphate and g/l with respect to sulfiz zgfi ig ig g gg lg i gg f iz g i i was P sg g 22 i p 3 A process according to claim l ?Nh ei'in t li e 22nis so utlon was treate wit m so 9 w wsulphuric acid- The heat of mixing caused some of the clzelntratlon of solids in the leach solution is from 1:4 to quor to evaporate and the final solution containing 890 5 g/l of sulphuric acid. The precipitated zinc sulphate A process which comprises m combmauon' a. roasting a zinc bearing material selected from the tsngngohydrate was filtered 0E and drymg w group consisting of iron-containing zinc sulphide b. leaching said calcine with water or an aqueous solution at a temperature in the range of 10C to 100C, using a solids: liquid ratio up to 2:1 to yield a leach slurry;

c. subjecting said leach slurry to a solids/liquid separation process to yield a leach solution which comprises as dissolved zinc sulphate at least 70 percent of the zinc contained in said calcine, and a residue, which contains silicaceous gangue, and any insoluble iron and lead compounds;

d. electrolysing said leach solution to yield, as a product, metallic zinc and, as an intermediate, spent electrolyte liquor, which is an aqueous solution of sulphuric acid and zinc sulphate;

e. concentrating sulphuric acid in said spent electrolyte liquor so that its liquid phase contains at least 600 g/l of H 80, and the bulk of the zinc sulphate is precipitated; and

f. separating this precipitate from the liquor and so obtaining sulphuric acid as a product, the precipitated zinc sulphate being redissolved in the free liquor to the electrolysis step and recycled in subsequent operations,

said zinc bearing material being roasted to step (a) in teo steps, wherein the first step comprises reasting said material in an atmosphere comprising oxygen and sulphur dioxide at a temperature in the range from 500 compounds. 

2. A process according to claim 1 wherein the spent electrolyte liquor is concentrated so that its liquid phase contains from 900 to 1,150 g/l of sulphuric acid.
 3. A process according to claim 1 wherein the concentration of solids in the leach solution is from 1:4 to 1:1.
 4. A process which comprises in combination: a. roasting a zinc bearing material selected from the group consisting of iron-containing zinc sulphide ore and iron-containing residues from a zinc manufacturing process at a temperature in the range 500*C to 800*C in an atmosphere which contains oxygen and sulphur dioxode to convert the zinc bearing material to a calcine comprising the bulk of the zinc originally present in the zinc bearing material in the form of zinc sulphate; b. leaching said calcine with water or an aqueous solution at a temperature in the range of 10*C to 100*C, using a solids: liquid ratio up to 2:1 to yield a leach slurry; c. subjecting said leach slurry to a solids/liquid separation process to yield a leach solution which comprises as dissolved zinc sulphate at least 70 percent of the zinc contained in said calcine, and a residue, which contains silicaceous gangue, and any insoluble iron and lead compounds; d. electrolysing said leach solution to yield, as a product, metallic zinc and, as an intermediate, spent electrolyte liquor, which is an aqueous solution of sulphuric acid and zinc sulphate; e. concentrating sulphuric acid in said spent electrolyte liquor so that its liquid phase contains at least 600 g/l of H2SO4 and the bulk of the zinc sulphate is precipitated; and f. separating this precipitate from the liquor and so obtaining sulphuric acid as a product, the precipitated zinc sulphate being redissolved in the free liquor to the electrolysis step and recycled in subsequent operations, said zinc bearing material being roasted to step (a) in teo steps, wherein the first step comprises reasting said material in an atmosphere comprising oxygen and sulphur dioxide at a temperature in the range from 500* to 800*C to convert the bulk of the zinc into zinc sulphate and the second step comprises maintaining the roast at a temperature in the range from 400* to 700*C in an atmosphere from which sulphur dioxide has been substantially purged to decompose water-soluble iron compounds. 